In modern pyrometallurgical practice sulphide concentrates are treated by using the difference in chemical affinity of metals to oxygen and sulphur contained in the raw material. To increase this difference, the smelting process is carried out in the presence of silicon dioxide. The silicate melt is then reduced. It should contain basic oxides required to ensure high rate of reduction. Thus, in this case use is made of the difference between elements in their affinity to oxygen.
Further progress in pyrometallurgical production of heavy nonferrous metals depends on the development of effective techiques for their recovery on the basis of autogenous smelting of sulphide material. Advantages of autogenous processes are known to permit high production rate (short time of holding the material in the oxidation zone); a sharp decrease in the amount of process gases; utilizing heating capacity of the concentrates and thus substantially diminishing the use of external heat sources; the possibility of effective treatment of the raw material which is relatively poor in nonferrous metals. In general, there is known a wide variety of autogenous processes. However, a feature most common to most of them is the use of highly developed surface of the sulphide material with a view to ensuring autogenous nature of the roasting and smelting processes.
For example, the Ontokumpu company, Finland, has developed a method of processing sulphide copper concentrates. The method in question is carried out by using a flow of possibly preheated air (that can be oxygen-enriched) for flash smelting of finely divided sulphide copper concentrates. The smelting process is effected in the presence of substantially silicate fluxes with the resultant dispersed molten mixture of silicate slag and matte containing not more than 65 percent by weight of copper.
This dispersed mixture of slag and matte is separated in accordance with specific weights. The matte is further treated until metallic copper is recovered therefrom, while silicate slag is depleted in an electric furnace by settling or flotation after being crushed and divided.
The gases resultant from flash smelting of the initial material contain from 2 to 20 percent of sulphurous anhydride. The loss of partially oxidized powder concentrate during smelting is in the range of 8 to 10 percent by mass of the material fed for smelting (see, for example, Shein Y. P. "Nonferrous Metals", No. 8, 1980, pp. 25-29 [Symposium on Nonferrous Metallurgy in Finland]; Engineering and Mining Journal, 1973, 174, No. 11, p. 103-108; M. J. Ethem "Erzemetall", No. 4, 29, 1976, pp. 182-186).
There exist other modifications of this method, which are directed at enhancing its technical-and-economic characteristics. For instance, the content of copper in the matte is brought up by increasing the concentration of oxygen in the flow of air supplied for smelting, the temperature of air or that of oxygen-enriched air being raised. It has been attempted to increase the rate of recovery of copper from raw material by arranging electrodes in the zone where slags settled down to become free from matte. However, the above-mentioned improvements failed to obviate serious disadvantages inherent in the prior-art method, namely:
the impossibility of processing raw material with higher-than-average content of zinc, or combining the processing of copper-zinc concentrates with the stage of producing blister copper;
a low rate of recovery of copper and zinc from silicate slags.
To overcome difficulties that might possibly occur in the process of producing rich matte (with over 65% by weight of copper), blister copper, or during treatment of a low-grade material containing difficulty oxidizable zinc sulphide, and in order to bring down the amount of magnetite (and the content of copper in slag), it has been attempted to distribute the supply of oxygen required for oxidation of sulphide sulphur, and to perform smelting along with preroasting and subsequent utilisation of fuel.
According to Pat. No. 2,515,464 (Federal Republic of Germany, cl. C 22 B 15/04) the initial material is first partially roasted and then smelted to produce matte. As this happens, a flow of hot oxygen is introduced into the roast gases which are then used as auxiliary air for combustion of fuel in the course of smelting. However, the above method is rather difficult to perform apart from being ineffective to prevent the formation of magnetite. In addition, it fails to solve the problems associated with the processing of copper-zinc or copper zinc-bearing concentrates and with the depletion of slags to permit good recovery of nonferrous metals therefrom.
There is known a KIVCET process which comprises roasting and smelting of flotation concentrates, containing nonferrous metals, in the atmosphere of commercial oxygen mixed with resusable dust resultant from cleaning of flue gases, and substantially by with silicate fluxes thereby obtaining matte and reducing zinc from the silicate melt in an electric furnace (see, for example, Japanese Pat. No. 16362/76, cl. C 22 B 15/00; V. V. Vylegzhanin et al. "Nonferrous Metals", No. 1, 1976, pp. 26-28; I. M. Cherednik et al. "Nonferrous Metals", No. 7, 1974, pp. 24-27; Melcher G. et al. Erzmetall, 1975, 28, No. 7/8, p. 313-322, I, II, III).
The KIVCET process is as follows.
A finely divided copper- or copper-zinc concentrate is submitted to flash smelting in the presence of substantially silicate fluxes in the atmosphere of commercial oxygen, followed by the formation of a dispersed mixture of silicate slag and copper matte. The dispersed melt is further separated in accordance with specific weights into silicate slag and matte. Next, the slag and matte are admitted into an electothermic furnace, wherein a solid carbonaceous material (coke breeze) is charged onto the surface of the molten silicate slag. Under the influence of high temperature the zinc oxide contained in the slag is reduced to metal which, when evaporized, passes into a vapour-gas phase while copper oxides, also present in the slag, are reduced to metal and metallic copper is settled down to form matte. The vapour-gas mixture is removed from the electric furnace, the vapour of metallic zinc is oxidized therein to form zinc oxide by means of air supply. The resultant zinc oxide is collected to be fed for further treatment. The matte is tapped from the electric furnace for further treatment to be blown through and oxidized with oxygen. As a result of this treatment, metallic copper is produced to undergo subsequent refining.
In the KIVCET process, due to the use of commercial oxygen for oxidizing finely divided sulphide material with a high specific surface, and owing to smelting carried out to produce matte (up to 55% by weight of copper and about 20% by weight of sulphur), it takes only about 0.1 s for the oxygen fed for smelting to be almost completely assimilated by the sulphide material in suspension. As a result, a high temperature is developed during smelting to ensure sufficiently high rate of oxidation of zinc sulphide to its oxide with zinc oxide passing into silicate slag.
However, the above method of processing sulphide copper- or sulphide copper-zinc concentrates with the production of silicate slags also suffers from a number of serious disadvantages, namely:
1. The specific production rate at the stage of slag depletion is low by reason of the nature of slag used (from 4 to 12 kg of zinc per day per sq. meter of area in the electric furnace); hence are appreciably high losses of heat (power input) for slag depletion, as well as labour input per unit of commercial product.
2. The degree of recovery of zinc from raw material to produce zinc oxide is not higher than 80 percent.
3. The impossibility to treat sulphide concentrates with a high content of zinc (about 20% by weight of zinc) because of a low rate and degree of oxidation of zinc sulphide and its conversion into slag so as to enable further reduction of zinc oxide from slag with a satisfactory rate of zinc recovery from the initial material (more than 60-70%).
4. The processing of the resultant copper matte to produce metallic copper requires substantial capital investment and labour input apart from the necessity to develop a complicated emission gas control system.
5. The impossibility to prevent the formation of considerable amounts of magnetite during flash smelting of concentrates; therefore, complete oxidation of the afore-mentioned sulphide material leeds to the formation of high-viscosity dispersed mixture of silicate slag, metallic copper or white matte, which is rather difficult to disintegrate. In addition, the recovery of copper and zinc from high-viscosity silicate slag is fairly intricate procedure hindering comprenensive treatment of low-grade and, in particular, copper-zinc concentrates.
From the above it follows that new methods are required to substantially increase the rate of production at the stage of slag depletion along with corresponding decrease in the labour and power inputs; to bring up the rate of zinc recovery from the initial material; to raise the rate of oxidation of sulphide concentrates during flash smetlting, thereby permitting the production of metallic copper to be effected directly at this stage, as well as to bring down capital investments and labour force involved in the production of metallic copper; to prevent the formation of considerable amounts of magnetite during flash smelting of sulphide concentrates, which will make it possible to introduce a wider variety of initial materials into the copper and zinc industry by making use of low-grade copper and copper-zinc sulphide concentrates along with direct production of metallic copper.
It is an object of the present invention to provide such a method of processing sulphide copper- or sulphide copper-zinc concentrates or mixtures thereof that will make it possible to enhance the production capacity and the rate of smelting of the initial material.
Another object of the invention is to provide such a method of processing sulphide copper- or sulphide copper-zinc concentrates or mixtures thereof that will make it possible to increase the production capacity at the stage of slag depletion among with a higher rate of zinc recovery from the initial material, while permitting production of metallic copper or white matte.
The foregoing objects are attained in a method of processing sulphide copper- or sulphide copper-zinc iron-bearing concentrates, which comprises flash smelting of said concentrates in the presence of fluxes selected from the group consisting of basic fluxes and combinations thereof with silicate fluxes, and oxygen with the resultant formation of a dispersed mixture of slag, metallic copper or white matte, subsequent reduction of copper and zinc oxides, contained in the molten slag with a solid carbonaceous material, followed by the formation of a vapour-gas mixture containing zinc vapour, and of metallic copper and slag poor in nonferrous metals, subjecting to oxidation the vapour-gas mixture containing zinc vapour, and collecting the resultant zinc oxide, wherein, according to the invention, the initial concentrates and fluxes are fed for smelting in amounts sufficient to ensure the production of a highly basic molten slag which, on being reduced to give off copper and zinc oxides, contains not more than 18% by weight of silicon dioxide. Owing to such conditions, calcium oxide forms low-melting oxysulphide eutectics with zinc and iron sulphides at the intermediate stages of oxidation. With high-melting and difficulty oxidizable zinc sulphide passing into a low-melting liquid phase (just as the product of its oxidation-zinc oxide), the rate of the charge desulphurization (the rate of oxidation) is markedly increased. Further, in the process of smelting carried out to produce highly basic slag with the afore-mentioned ratio of the basic slag-forming components, there takes no place or hardly takes place any formation of magnetite which otherwise greatly increases viscosity of conventional silicate slags and thus hinders the process of oxidation. The present invention permits low-melting calcium ferrite to be formed instead of magnetite so as to dilute the products of oxidation and in this way to ensure the transportation of oxygen in the volume of droplets of the dispersed oxysulphide melt. Owing to these two main factors, that is--the formation of low-melting oxysulphide phases by means of calcium oxide and zinc and iron sulphides, and the formation of low-melting calcium ferrite instead of magnetite, it became possible to substantially intensify the process of desulphurization so that the copper sulphides present in copper are passed into metallic copper directly at the stage of oxidizing-flash smelting. In other words, copper is now produced directly from sulphide copper- or sulphide copper-zinc concentrates with the effect that the conversion of matte is ruled out to permit substantial cut-down of expenses and labour involved. Moreover, a considerable increase in the rate of oxidation of zinc sulphide has made it possible to perform effective treatment of zinc-bearing sulphide concentrates. This, in turn, greatly expands the variety of initial materials in the copper and zinc industries by making use of a low-grade sulphide material from which valuable components (copper, zinc and sulphur) are successfully extracted. Furthermore, since highly basic slag has a lower viscosity than silicate slag, the metallic copper or white matte formed during smelting is easily separated from the highly basic molten slag and passes into the bottom phase. This brings down the amount of copper lost with slag in the form of mechanical suspension (about 10 rel.%). The highly basic molten slag obtained during flash smelting contains the whole of zinc, a substantial amount of lead, with the content of copper in the form of oxide constituting about 5% by weight. The reduction of oxides of nonferrous metals (copper, zinc, lead) from highly basic molten slags proceeds at much higher rate than in the case of silicate slags and approaches the specific rate of the fuming process. In part, this takes place due to a higher intensity of the following reaction: EQU (n+2m)MeO+(n+m)C=(n+2m)Me+nCO+mCO.sub.2 ( 1)
with a corresponding increase in the activity of oxides of nonferrous metals contained in highly basic slags. Another reason for this are key changes in the coordination structure of metals in the highly basic slags as compared to silicate slags. In highly basic slags, the structure-determining component is calcium (coordination number--six) instead of silicon in the silicate slags (coordination number--four). Such structural transformations lead to a decrease in the amount of free energy and especially in that of free energy resultant from the activation of the following reaction: EQU FeO.sub.s1 =Fe.sub.s1 +Fe.sub.2 O.sub.3 s1 ( 2)
equivalent to the increase of its velocity by several orders. As a result, zinc, copper and lead oxides are reduced at a much higher rate in accordance with the following reactions: EQU ZnO+Fe.sub.s1 .fwdarw.Zn.uparw.+FeO (3) EQU Cu.sub.2 O+Fe.sub.s1 .fwdarw.2Cu+FeO (4)
Thus, during carbon reduction of nonferrous metals from highly basic slags it is more difficult, as compared to silicate slags, for metallic iron to be separated into an independent phase or into blister copper. Therefore, with the content of nonferrous metals in the reduced slag amounting to 1-1.5% by weight (0.5 to 0.7% by weight of zinc, 0.3 to 0.5% by weight of copper, 0.01% by weight of lead), metallic copper containing 1 to 3% by weight of iron is found in equilibrium therewith. In contrast, the content of zinc in silicate slags under similar conditions of the melt reduction by means of solid carbonaceous material can not be lower than 3.5 to 4% by weight of zinc, which is due to the fact that substantial amount of iron passes into the bottom phase with its melting temperature sharply raising to render impossible deep depletion of silicate slag because of solidification of the bottom phase. Thus, for the total content of nonferrous metals to be in the range of 1-1.5% by weight, it takes 10 to 15 times more time for the depletion of silicate slags than for the depletion of highly basic molten slags. Furthermore, such deep depletion results in that copper is considerably contaminated with metallic iron (not less than 20% by weight of iron) with the melting temperature thereof being over 1400.degree. C. From the above it follows that in comparison with the Japanese Pat. No. 16362/76 cl. C 22 B 15/00, it becomes possible not only to improve the process characteristics at the stage of flash smelting of sulphide concentrates to produce highly basic slags in the presence of oxygen, but to improve the process characteristics at the stage of slag depletion by using a solid carbonaceous material. To develop the point, the amount of zinc recovered from the initial material is substantially increased (from 75-80% to 94-97%), the rate of sublimation of volatile zinc is enhanced (from 4-12 to 40-60 kg/m.sup.2 per day), with power and labour inputs per unit of commercial product (zinc oxide) being proportionally reduced.
According to the invention, the content of silicon dioxide in the depleted highly basic molten slag is preferably maintained within the range of 3 to 16% by weight with the use of silicate flux, whereafter the melt is cooled down to a temperature of its complete solidification at a rate of 0.5 to 60 degrees per minute with the resultant production of self-disintegrating material from which nonferrous metals are finally recovered. On the one hand, this will ensure moderate melting temperature of the melt, and the formation of sufficient amounts of dicalcium silicate during cooling and solidification of the melt, on the other. As the depleted highly basic molten slag is cooled down to a temperature of its complete solidification at a rate of 0.5 to 60 degrees per minute, this silicate is produced in the form of fairly pure large-size crystals (with average diameter of 40-80 mcm). At a temperature of about 675.degree. C., the crystals of dicalcium silicate undergo polymorphic transformation, which is followed by a 12 to 15% increase in their volume. This, in turn, brings about interior strain in the slag monolith. Under the influence of this interior strain the monolith is spontaneously disintegrated to form particles of a size fit for flotation (the yield of this fraction is 75-85%). In addition, the temperature decrease at the above-indicated rate makes possible the separation of copper left in the slag in the form of droplets with a diameter of 80 mcm (about 70% of copper is recovered from the slag). Due to spontaneous disintegration of slag and with separation of copper therefrom in the form of droplets, the surfaces of which become exposed during spontaneous disintegration of the slag monolith, final extraction of copper from the depleted highly basic slag is simplified. Thus, with the content of copper in the depleted slag of 0.3 to 0.5% by weight, it is possible to obtain a concentrate by means of flotation with the content of copper ranging from 10 to 15% and with the discarded material containing from 0.1 to 0.15% by weight of copper (depending upon the initial content of copper). At the same time, following depletion, comminution and treatment of silicate slags by flotation, they still contain, as a rule, 0.5% by weight of copper (at best 0.3% by weight of copper). Therefore, when smelting is carried out to produce highly basic slags to undergo further treatment with carbonaceous material in order to obtain self-disintegrating depleted slag, the most arduous operations of the separating stage, such as crushing and dividing the solid monolith, are ruled out; the loss of copper with the discarded material is reduced from 1-1.5 rel.% to 0.3-0.6 rel.% (in contrast to the smelting followed by the formation of silicate slag).
According to the invention, the resultant metallic copper is preferably submitted to refining in the presence of silicate fluxes for the purpose of producing refined copper and silicate slag. It has been found that the metallic copper produced during treatment of copper- or copper-zinc sulphide concentrates contains impurities of lead, zinc, iron and other elements to be removed from metallic copper. As molten metallic copper is blown through with air in the presence of silicate fluxes, the above-mentioned impurities are ozidized to pass into silicate slag where oxides of the metallic impurities are combined to form strong low-volatile silicates and thus are removed from metallic copper. Simultaneously, copper is partially ozidized to form copper oxide which also passes into silicate slag and is partially dissolved in metallic copper to oxidize metallic impurities. Since during refining of metallic copper the melt is intensively mixed with air, the dissolving of silicate flux proceeds at a rapid rate even when large-size lumps of silicate flux are used. Therefore, with a view to reducing the loss of heat during transportation of metallic copper for refining, the latter is preferably delivered to this stage in molten state.
According to the invention, the metallic copper obtained during flash smelting of concentrates in the presence of oxygen in the course of treatment of the dispersed mixture of highly basic molten slag, metallic copper or white matte with a solid carbonaceous material, and the metallic copper obtained during treatment of highly basic molten slag with a solid carbonaceous material, are discharged separately.
It has been found that during flash smelting of copper- or copper-zinc concentrate in the presence of basic fluxes, the resultant metallic copper contains up to 90% of copper contained in the initial material fed for smelting. The metallic copper produced at the stage of flash smelting is preferably withdrawn from the process with only highly basic oxide melt being fed for carbon treatment. This permits the bulk of copper (up to 90 rel.%) to be obtained in the form of commercial product, the purest possible for the given type of initial material, whereas the copper residue in the oxide melt is preferably extracted in the form of less pure metal at the stage of carbon treatment of slag. The foreign-metal impurities such as lead, zinc and iron pass into metallic copper mostly in the form of metals which are obtained as intermediate products resultant from the reaction between corresponding sulphides and oxides, and according to reactions (2) and (3). Since in the process of flash smelting, due to high oxidation potential of the medium, the concentration of such intermediate products cannot be high (it has maximum value for lead), as also the degree of transition of lead, zinc and iron into metallic copper. At the stage of carbon treatment of the highly basic melt, the oxides of copper, lead, zinc and, in part, of iron are reduced to metals. It should be observed that with the residual total content of copper, zinc and lead in the oxide melt being 1 to 1.5% by weight, metallic copper containing up to 8% by weight of lead, up to 1.5% by weight of zinc and up to 2% by weight of iron will be in equilibrium therewith. However, since the bulk of copper contained in the initial material is recovered prior to the stage of carbon treatment in the purest possible form, the overall contamination of metallic copper with the above-mentioned impurities turns out to be 2 to 3 times lower during separate tapping in contrast to the practice when all copper is tapped following the stage of slag treatment. In addition, during separate discharging of metallic copper, direct recovery of zinc and lead in the form of sublimates inside the furnace is increased by 6 to 8 rel.%, which brings down the losses of these metals with metallic copper.
According to the invention, the silicate slag produced during refining of metallic copper is preferably used as the silicate flux, which may contain up to 50% by weight of copper, up to 10% by weight of zinc, up to 10% by weight of lead, since these metals are present in the slag basically in the form of oxides, the slag should be subjected to treatment with carbonaceous material. On the other hand, the slag resultant from refining of metallic copper contains silicon dioxide which is preferably introduced for processing copper- and copper-zinc sulphide concentrates so as to obtain a self-disintegrating slag monolith required for subsequent final recovery of nonferrous metals. Both of these tasks are successfully fulfilled if the above-mentioned slag is used as the silicate flux.
According to the invention, the silicate flux is preferably introduced for use in flash smelting of the concentrate. Due to the fact that high temperature (over 1500.degree. C.) is developed in the course of flash smelting of sulphide concentrate in the presence of oxygen, a refractory material, such as quartz sand reduced to a size of minus 0.5 mm, is preferably used as the silicate flux. The presence of silicon dioxide in the furnace burden brings down the effect of formation of oxysulphide phases on the base of calcium oxide and copper and zinc sulphides. However, since silicate flux is fed for smelting in the form of a refractory material, the dissolving of silicon dioxide in the resultant highly basic melt takes place mainly when the process of oxidation of metal sulphides in suspension has been completed. Therefore, adversary effect of silicon dioxide on the flash smelting of sulphide concentrate turns out to be insignificant, with refractory silicate material being successfully used in the process.
According to the present invention, the silicate flux is preferably introduced into the molten highly basic slag produced flash smelting of the concentrate. This will make it possible to reduce the load with regard to the inert material (silicon dioxide) at the stage of flash smelting of sulphide concentrate. At the same time, this allows for the heat liberation processes during oxidation of metal sulphides and the heat absorption processes during dissolution of silicate flux in the highly basic melt to be run at minimum intervals. This technique is advised when the silicate slag resultant from refining metallic copper is used as the silicate flux, which containts all components predominantly in oxidized state. Since this type of silicate slag has relatively low temperature (about 1200.degree. C.), it is readily dissolved in the dispersed highly basic slag having a temperature of more than 1500.degree. C. However, to avoid oversaturation of the resultant melt with silicon dioxide and settling of refractory dicalcium silicate, the silicate slag from refining of metallic copper should be fed at regular intervals.
According to the invention the silicate flux is preferably introduced into the depleted highly basic molten slag, which permits proportional reduction in load with regard to an inert material at the stage of flash smelting of sulphide concentrate and at the stage of treatment of the highly basic molten slag with a solid carbonaceous material. In addition, the presence of silicon dioxide at the stage of treatment of the highly basic melt somewhat brings down the process rate and leads to an increase in the melting temperature of the melt (up to 1330.degree. C., with the content of silicon dioxide in the melt being 20% by weight). Therefore, when treatment of the highly basic melt is carried out with the content of silicon dioxide (determined by the composition of sulphide concentrate) being at its lowest, the rate of flash smelting for the given type of concentrate is the highest and the rate of recovery of nonferrous metals from the highly basic slag is maximal.
According to the present invention, the depleted silicate slag poor in nonferrous metals is preferably used as the silicate flux. This is advised when silicate flux is introduced into the highly basic slag poor in nonferrous metals. However, the content of nonferrous metals (copper, lead) should be more than 0.5% by weight. Such a low content of nonferrous metals in silicate slag permits their extraction to produce the concentrate after obtaining self-disintegrating slag monolith and final recovery of nonferrous metals, for example, by flotation methods. The current pyrometallurgical techniques used for smelting copper with the resultant formation of silicate slag make it possible to bring the content of copper therein up to 0.5% by weight, this being much higher than with the method of the present invention (0.1-0.15% by weight of copper). Therefore, such depleted silicate slags may be used in the given method.
Further objects and advantages of the present invention will become more apparent to those skilled in the art upon a further reading of this disclosure, particularly when viewed in the light of illustrative examples.